Recovery of aluminium and fluoride values from spent pot lining

ABSTRACT

Process for the recovery of aluminum and fluoride values from spent pot lining materials comprising the steps of calcining spent pot lining material to produce an ash having environmentally acceptable levels of cyanide contamination, subjecting the ash to a leaching step in a solution containing a mineral acid and a corresponding aluminum salt in such proportions as to dissolve the aluminum and fluoride values, and subjecting the leached liquid to thermal hydrolysis to cause precipitation of an aluminum fluoride product.

FIELD OF INVENTION

This invention relates to the recovery of fluorine and aluminum valuesfrom waste materials, and more particularly to the recovery of aluminiumfluoride from spent pot lining materials obtained from electrolyticreduction cells used to produce aluminium metal.

BACKGROUND OF THE INVENTION

The carbon lining which forms the internal, side and bottom walls of anelectrolytic reduction cell gradually degrades due to the extremely hightemperatures and the corrosive conditions that exist during operation ofthe cell. This degradation gradually causes failure of the carbon blockswhich make up the cell, allowing molten aluminium to penetrate thecarbon blocks which often causes distortion of the cell. At this stagethe cell is removed from service in the pot line. In addition to thecarbonaceous cathode material, the refractory and insulating materialsthat surround the cell as well as the steel cathode bars imbedded in thebottom of the cell are also removed. This material is called spent potlining or SPL.

During its life the carbon lining of the electrolysis cell absorbsquantities of the bath materials which include aluminium metal, sodiumaluminium fluorides and other fluorides. Aluminium carbides and nitridesare also formed during cell operation and these are also deposited inthe carbonaceous cathode material. Spent pot lining is currently listedas a hazardous waste by the U.S. Environmental Protection Agency as itcontains potentially harmful leachable cyanides and fluorides that canenter the ground water during open air storage. As well, ammonia,hydrogen, hydrogen cyanide, methane and phosphine are produced when thematerial becomes wet. Various disposal techniques where the SPL caneither be destroyed or the materials in the SPL used in other industrialprocesses have been developed over the years. However, of theseprocesses none has been accepted as standard practice in the industry.

These include processes where the spent pot lining can be used as areplacement for fluorspar in the steel industry or where the carbonvalues of the spent pot lining are used as a supplementary source offuel. For example spent pot lining has been burnt in cement kilns andhere the cyanide is destroyed while both the carbonaceous material andthe fluoride values are used.

Although fluidized bed combustion techniques for disposing a spent potlining has been demonstrated on a pilot scale, this technique has yet tobe demonstrated on a commercial scale. While the cyanide levels of thespent pot lining have been reduced to acceptable levels in suchinstances, the ash still contains the fluorides which have to beimmobilized for example with calcium hydroxide, if the ash is to bedisposed as landfill.

As the spent pot lining contains significant amounts of fluorinecontaining chemicals as well as appreciable amounts of aluminium whichcan be recycled into the aluminium smelting process, there is aneconomic incentive both for recovering these values and for producing aspent pot lining residue which can be disposed in an environmentallyacceptable manner.

Examples of processes aimed at recovering aluminium fluoride from spentpot lining materials are to be found in U.S. Pat. Nos. 4,508,689 Bush etal, 4,597,953 Bush and 4,889,695 Bush. In these processes, the potlining material is crushed and leached to extract the fluoride andaluminium values. However, the processes described in the above patentsdo not address the problem of disposal of the potentially dangerouscyanide containing residue and the fluoride and aluminium value recoveryrate is not particularly high.

A membrane process for treating SPL has been developed which uses lowtemperature solution processing using the well known cryolite recoverytechnology. However, the solution still contains silica, aluminium, ironand cyanide and these components can lead to membrane fouling.

SUMMARY OF INVENTION AND OBJECT

It is an object of the present invention to provide an improved processfor the recovery of aluminium and fluoride values from SPL in which thedifficulties associated with cyanide contaminated residues aresignificantly reduced.

The invention provides a process for the recovery of aluminium andfluoride values from spent pot lining materials, comprising the steps ofcalcining spent pot lining material to produce an ash havingenvironmentally acceptable levels of cyanide contamination, subjectingthe ash to a leaching step in a solution containing a mineral acid and acorresponding aluminium salt in such proportions as to dissolve thealuminium and fluoride values, and subjecting the leached liquid tothermal hydrolysis to cause precipitation of an aluminium fluorideproduct.

It will be appreciated that the process defined above produces arelatively high purity aluminium fluoride product without the use ofcaustic soda solutions, thereby reducing the amount of by-productliquors that need to be treated. Similarly, since the SPL ash which issubjected to the leaching step is substantially free of cyanidecontamination, the process according to the invention represents anenvironmentally acceptable process for the recovery of an aluminiumfluoride product from the spent pot lining material. In addition, thecalcining of the SPL not only substantially removes cyanidecontamination, it also significantly frees the aluminium and fluorinevalues for recovery by the chosen leaching process thereby maximisingthe desired recovery process.

In a preferred form of the invention, the SPL is first crushed to aparticle size less than 600 micrometers and the particles are subjectedto calcination in a furnace operating at a temperature in the range 680°C. to 850° C. to produce a low-cyanide ash. The ash is then added to asolution containing a mineral acid comprising hydrochloric, sulfuric ornitric, or mixtures thereof, and the equivalent aluminium salt, i.e.aluminium chloride, aluminium sulphate or aluminium nitrate. The amountof water used is such that the ratio of water to SPL ash is in the range5 to 25. The amount of aluminium salt used is such that the salt to SPLash ratio is in the range 0.1 to 0.8 while the ratio of acid to SPL ashis in the range 0.1 to 1.2. The SPL ash is leached with agitation attemperatures between 40° and 100° C. The time of leaching being between5 and 360 minutes. In some instances it may be preferable to add thealuminium salt after the SPL has been partially leached with the acidonly. Another method of adding the aluminium salt to get the above saltto SPL ash ratio is to produce it in-situ by reacting an aluminiumcompound e.g. aluminium hydroxide with the appropriate acid.

An alternative and preferred ash dissolution method, to prevent silicadissolution is as follows. The spent pot lining ash is first mixed withconcentrated sulfuric acid. Water is then added to the acid/SPL ashmixture and the resultant mixture is allowed to age at temperaturesbetween ambient and 150° C. for times up to 24 hours. The aged acid/SPLmixture is then leached with an aluminium sulphate solution to producean Al₂ (SO₄)₃ /H₂ SO₄ ratio in the range 0.75 to 1 and at temperaturesup to 100° C. and for times up to 3 hours. The leach liquor is thenfiltered from the residue which contains mainly carbon and silicatematerial.

The filtrates from the acid leaching processes, which contain aluminium,sodium, iron and minor amounts of calcium and magnesium as well as thefluoride and the anion of the leaching acid is then placed in anautoclave. The vapor space above the liquid is purged of air with aninert gas to exclude oxygen and the over-pressure above the liquid isadjusted with the inert gas in order to prevent precipitation of ironoxide. The pressure above the liquid before heating commences may bebetween atmospheric and 2000 kPa. Alternatively, the vapor space abovethe liquid may be evacuated to achieve similar results.

The contents of the autoclave are then heated with agitation attemperatures from about 105° C. with a corresponding pressure of about20 kPa to about 265° C. with a corresponding pressure of about 5000 kPa.Depending on the temperatures used the holding times vary from 1 minuteto over 5 hours. At the conclusion of the hydrothermal precipitation thereactor is cooled and the pressure reduced to atmospheric. The solutioncontains a white solid which is filtered from the barren liquor anddried. X-Ray Diffraction scans on the dried powder show it to beAl(OH,F), 0.375 H₂ O.

The alternative processes outlined above are shown schematically in FIG.4 of the drawings.

It will be appreciated that one of the major advantages of thispreferred processing method is the production of an aluminium fluorideproduct low in silica and iron. Since this method of processing is, forthis reason, also likely to be useful in the processing of uncalcinedpot lining materials, a further aspect of the invention provides: amethod of processing spent pot lining materials to produce a productcontaining aluminium and fluoride values which is low in iron andsilica, comprising mixing the spent pot lining material with aconcentrated mineral acid solution, ageing the mixture for a period ofup to about 24 hours to prevent dissolution of the silica contained inthe mixture, leaching the mixture by the addition of an aluminium saltcorresponding to the mineral acid in such proportions as to dissolve thealuminium and fluoride values, and subjecting the leached liquid tothermal hydrolysis in the presence of an inert atmosphere or vacuum toprevent precipitation of the iron in the leached liquid and to causeprecipitation of a product containing aluminium and fluoride values.

One significant disadvantage of the above method using uncalcined spentpot lining materials is the presence of cyanide contamination in theleached liquid, and for this reason it is preferred that the spent potlining materials should be calcined in the manner defined above beforeperforming the method last defined above. However, it may neverthelessbe sufficiently advantageous to have a precipitate substantially free ofiron and silica to warrant the separate treatment of the leached liquidfollowing precipitation to remove the cyanide contamination. In thisregard, it will be appreciated from a consideration of U.S. Pat. No.4,889,695 that a separate and distinct deironing step is usuallynecessary and that the process described in this patent does not addressthe question of silica contamination and its attendant disadvantages.

BRIEF DESCRIPTION OF THE DRAWINGS

The following examples illustrate the invention and are described withreference to the accompanying drawings in which:

FIG. 1 is a Sankey diagram showing the mass balance of a systemembodying the invention;

FIG. 2 is a graph showing the total cyanide content of the productsstreams throughout the test;

FIG. 3 is a graph showing the carbon content of the feed and productduring the test, and

FIG. 4 is a schematic flow diagram illustrating the steps involved inthe alternative processes embodying the invention.

The invention is more clearly illustrated by the following exampleswhich are intended to be illustrative only and should not be taken aslimiting the invention in any manner. The SPL ash referred to in thefollowing examples was prepared in a fluidized bed contactor of the typedescribed in our copending International Patent ApplicationPCT/AU91/00342 in the name of Comalco Aluminium Limited, the contents ofwhich are incorporated hereby by cross-reference. The crushed spent potlining material, having a particle size of less than 600 micrometers,was subjected to calcining temperatures in the range 680° C. to 850° C.,and most typically about 720° C.

Test Details

The feed material was less than 600 micrometers low carbon SPL. Atypical composition is shown in Table 1.

                  TABLE 1                                                         ______________________________________                                        TYPICAL -600 MICROMETERS                                                      LOW CARBON FEED COMPOSITION                                                   % wt                                                                          C      Al        Si    Na      F    ppm CN                                    ______________________________________                                        16.9   22.5      7.8   11.7    13.3 200-500                                   ______________________________________                                         The tests were conducted under the following conditions:                      Bed Temperature (Target) = 720° C.                                     Fluidizing air rate = 185 kg/hr                                               Chamber pressure = -40 mm H.sub.2 O (gauge)                                   Feedrate = 28 kg/hr                                                      

During the tests, samples of product ash were collected hourly forchemical and screen analysis. Gaseous cyanide fluoride and sodiumemissions were sampled from the exhaust duct. The mass flowrate of allsolid streams was measured hourly and all relevant variables were loggedcontinuously from the microprocessor onto magnetic disk.

RESULTS AND DISCUSSION System Performance

The two 5-hour tests proceeded smoothly with no operational problems. Noagglomeration was encountered in the system. The mass balance ispresented in Sankey diagram form in FIG. 1.

Total Cyanide Content

FIG. 2 shows the total cyanide content of the product streams throughoutthe test. The cyanide content of a typical low carbon feed is shown forcomparison. The reduction in total solid cyanide is 80%-from 340 ppm inthe feed to 70 ppm in the combined product.

Carbon Burn-out

The carbon content of feed and product throughout the test is shown inFIG. 3. Average carbon burn-out was 58%.

Gaseous Emissions

During each of the 5-hour test periods, 2 samples of off-gas were takenand independently analysed for gaseous F⁻ CN⁻ and Na. The full NATAreport is summarized in Table 2.

                  TABLE 2                                                         ______________________________________                                        OFF-GAS ANALYSIS                                                              Sam- T(°C.)                                                                         Mass                                                             ple  at      flow   F           Mass   CN                                     Pe-  Sam-    (mg/   Concn.                                                                              %     flow   Concn.                                                                              %                                riod pling   hr)    (ppm) Volat.                                                                              (mg/hr)                                                                              (ppm) Volat.                           ______________________________________                                         1*  536     960    5.82  <0.1  55.8   0.27  0.6                              2    552     90     0.55  <0.01 29.4   0.14  0.3                              3    561     66     0.44  <0.01 31.2   0.16  0.3                              4    558     72     0.45  <0.01  30.06 0.15  0.3                              ______________________________________                                         *Particulate breakthrough occurred during sampling                            Examples of the leaching and pressure precipitation steps defined above       are as follows:                                                          

EXAMPLE I

Thirty grams of SPL ash was mixed with 12.0 grams of concentratedsulfuric acid. When this was well mixed 3.7 grams of water was added andthe mixture was allowed to age at room temperature for 2 hours. The agedmixture was added to 470.0 grams of water to which 35.0 grams of Al₂(SO₄)₃, 18H₂ O had previously been added. The temperature of thesolution was gradually increased to 93° C. over an hour. The solutionwas allowed to cool overnight and was filtered the next morning. X-RayFluorescent analysis of the dried leach residue indicated that 97% ofthe fluorine, 47% of the aluminium, 69% of the iron and 96% of thesodium had been extracted from the SPL ash. ICP analysis of the solutionshowed that it contained 84.1 g/l SO₄, 0.7 g/l Fe, 18.2 g/l Al, 16.4 g/lNa and 1.0 g/l SiO₂.

EXAMPLE II

A second 30.0 grams of SPL ash was again mixed with 12.0 grams ofconcentrated sulfuric acid but in this case 5.5 grams of water was addedto the mixture. The mixture was again aged at room temperature for 2hours. The aged mixture was added to 470.0 grams of water containing35.0 grams of Al₂ (SO₄)₃, 18H₂ O which was heated to 80° C. The solutiontemperature was gradually increased to 91° C. over an hour. The solutioncooled overnight and was filtered the next morning. X-Ray Fluorescentanalysis of the dried leach residue showed that 98% of the fluorine, 47%of the aluminium, 72% of the iron and 96% of the sodium had beenextracted from the SPL ash. In this case the solution contained 86.6 g/LSO₄, 0.7 g/l Fe, 19.1 g/l Al, 17.0 g/l Na and 0.9 g/l SiO₂.

EXAMPLE III

A solution was prepared which contained 305.96 g of the leach solutionfrom example I and 965.80 g of the leach solution from example II. Thiswas added to an autoclave and heated to 200° C. and held at thistemperature for 3 hours. During the heating time the solution wasagitated at 500 rpm. When the autoclave had cooled and was opened thesolution contained a pale pink solid. Upon filtering and drying 28.55grams of a pink solid was obtained, X-Ray diffraction of the dried solidshowed that it contained Al(OH,F)₃ 0.375 H₂ O and Fe₂ O₃. X-RayFluorescent analysis of the dried solid showed that it contained 26.6%F, 24.8% Al, 0.21% Fe, 1.3% Na, 0.08% Mg and 0.16% Ca. The silicacontent was below 0.2%.

EXAMPLE IV

607 grams of a solution prepared in a similar manner to those inexamples I and II was added to the autoclave. Dry nitrogen was admittedfor five minutes to displace the air above the liquid. The nitrogenover-pressure was then increased so that the pressure above the liquidwas 210 kPa. The solution was then heated to 150° C. while it was beingagitated at 500 rpm. It was held at this temperature for 1 hour. Whenthe autoclave was cool the slurry was filtered and the solids dried.From this test 18.15 grams of white powder were obtained which XRDanalysis identified as Al(OH,F)₃, 0.375 H₂ O. The solids were analysedby XRF and gave the following analysis; 0.15% CaO, 0.05% Fe₂ O₃, 0.06%P₂ O₅, 47.41% Al₂ O₃, 0.21% MgO, 0.35% Na₂ O and 25.89% F. Silica and K₂O were below the detection limit of the equipment. Loss on Fusion of thesample was 26.32 %.

EXAMPLE V

607 grams of a solution prepared as for the previous example was addedto the autoclave. Dry nitrogen was used to purge the air from above theliquid for 5 minutes. The nitrogen over-pressure was then increased to210 kPa. The solution was heated to 200° C. while agitating at 300 rpmand held at this temperature for 63 minutes. When the autoclave was coolthe slurry was filtered and the solids dried. From this test 22.13 gramsof white powder was obtained. XRF analysis showed that it contained0.20% CaO, 0.05% Fe₂ O₃, 0.04% P₂ O₅, 49.12% Al₂ O₃, 0.22% MgO, 1.105Na₂ O and 25.99% F. Silica and K₂ O were again below the detection limitof the equipment. Loss on Fusion of this sample was 24.01%.

Tables 4 and 5 detail the temperature, pressure and time parameters offurther experimental tests conducted in accordance with the invention,while Table 4 details the sample analysis of five of the tests toillustrate the significantly reduced silica content of samples treatedby the preferred process 2 shown schematically in FIG. 4 of thedrawings.

                  TABLE 3                                                         ______________________________________                                        TEST PARAMETERS                                                                       Temp    Pressure   Time  Solids Produced                              Run     (°C.)                                                                          (kPa)      (min) (g)                                          ______________________________________                                        HP1     200     1600       61    33.2                                         HP2     200     1800       61    30.4                                         HP3     200     1400       186   30.4                                         HP4     200     1500       181   28.6                                         HP9*    200     2300       63    22.1                                         HP11*   200     2200       15    23.3                                         HP6     180      910       180   24.5                                         HP7     150      350       72    18.4                                         HP8*    150     1200       79    18.2                                         HP10*   120      910       60    8.2                                          HP13*   110      770       120   6.1                                          HP43    200     2100       40    30.6                                         HP44    200     2100       40    35.5                                         HP45    190     2100       40    29.3                                         ______________________________________                                         *Nitrogen over pressure of 210 kPa.                                      

                                      TABLE 4                                     __________________________________________________________________________    SAMPLE ANALYSIS                                                               Sample                                                                            CAO FE203                                                                             P205                                                                             AL203                                                                             SI02                                                                              K20 MGO NA20                                                                              F  LOF                                                                              LOI %                                __________________________________________________________________________    HP8 0.15                                                                              0.04                                                                              0.06                                                                             47.12                                                                             -0.03                                                                             -0.01                                                                             0.22                                                                              0.42                                                                              26.43                                                                            26.32                                                                            PPTE*                                HP9 0.20                                                                              0.05                                                                              0.04                                                                             48.92                                                                             -0.04                                                                             -0.00                                                                             0.21                                                                              1.08                                                                              24.31                                                                            24.01                                                                            PPTE*                                HP43                                                                              0.18                                                                              0.03                                                                              0.05                                                                             46.70                                                                              5.07                                                                             -0.01                                                                             0.16                                                                              0.82                                                                              26.44                                                                            21.62                                                                            PPTE+                                HP44                                                                              0.43                                                                              0.09                                                                              0.04                                                                             43.13                                                                              4.23                                                                             -0.00                                                                             0.14                                                                              2.40                                                                              23.58                                                                            20.82                                                                            PPTE+                                HP45                                                                              0.20                                                                              0.05                                                                              0.05                                                                             46.46                                                                              4.96                                                                             -0.01                                                                             0.14                                                                              0.78                                                                              24.45                                                                            22.26                                                                            PPTE+                                __________________________________________________________________________     *Process 2 (FIG. 4)                                                           +Process 1 (FIG. 4)                                                      

Although the process as outlined can produce a high purity aluminiumfluoride product without the use of caustic soda, in certaincircumstances it may be advantageous to increase the pH of the leachsolution as this may facilitate removal of any residual fluoride fromwaste solutions. This pH adjustment can be accomplished by the additionof caustic soda, ammonia, caustic magnesia or other basic compounds.

I claim:
 1. A process for the recovery of aluminium and fluoride valuesfrom spent pot lining materials, comprising the steps of calcining spentpot lining material to produce an ash having environmentally acceptablelevels of cyanide contamination, subjecting the ash to a leaching stepin a solution containing a mineral acid and a corresponding aluminiumsalt in such proportions as to dissolve the aluminium and fluoridevalues, and subjecting the leached liquid to thermal hydrolysis to causeprecipitation of an aluminium fluoride product.
 2. The process of claim1, wherein the spent pot lining material is first crushed to a particlesize less than 600 micrometers.
 3. The process of claim 1 or 2, whereinthe mineral acid is selected from hydrochloric, sulfuric or nitricacids, or mixtures thereof, and the corresponding aluminium saltcomprises aluminium chloride, aluminium sulphate or aluminium nitraterespectively, or mixtures thereof.
 4. The process of claim 3, whereinthe amount of aluminium salt used in such that the salt to SPL ash ratiofalls substantially in the range 0.1 to 0.8 while the ratio of acid toSPL ash falls substantially in the range 0.1 to 1.2.
 5. The process ofany preceding claim, wherein the solution contains water in a ratio toash which falls substantially in the range 5 to
 25. 6. The process ofany preceding claim, wherein the ash is leached with agitation at atemperature which substantially falls within the range 40° to 100° C.over a period which substantially falls within the range of 5 to 360minutes.
 7. The process of any preceding claim, wherein the aluminiumsalt is added after the ash has been partially leached with the mineralacid only.
 8. The process of any preceding claim, wherein the aluminiumsalt is produced in-situ by reacting an aluminium compound with themineral acid.
 9. A process for the recovery of spent pot liningmaterials to recover aluminium and fluoride values comprising mixing thespent pot lining material with a concentrated mineral acid solution,ageing the mixture for a period of up to about 24 hours to preventdissolution of the silica contained in the mixture, leaching the mixtureby the addition of an aluminium salt corresponding to the mineral acidin such proportions as to dissolve the aluminium and fluoride values,and subjecting the leached liquid to thermal hydrolysis in the presenceof an inert atmosphere or vacuum to prevent precipitation of the iron inthe leached liquid and to cause precipitation of a product containingaluminium and fluoride values.
 10. The process of claim 9, wherein thespent pot lining material is calcined to produce an ash havingenvironmentally acceptable levels of cyanide contamination before thefirst step in the process.
 11. A process for the recovery of aluminiumand fluoride values from spent pot lining materials, comprising thesteps of calcining, spent pot lining material to produce an ash havingenvironmentally acceptable levels of cyanide contamination, mixing theash with concentrated sulfuric acid, adding water and allowing theresultant solution to age at temperatures about falling in the range ofabout ambient to about 150° C. for a time period of up to twenty fourhours, leaching the aged solution with an aluminum sulphate solution toproduce a solution having an Al₂ (SO₄)₃ /H₂ SO₄ ratio falling within theof about range 0.75 to about 1 at temperatures up to about 100° C. for atime period up to about 3 hours, filtering the leached liquor from theresidue to separate a carbon and silicate residue, placing the filtratein an autoclave from which air is of about excluded, and adjusting theover-pressure within the autoclave to prevent precipitation of ironoxide, heating the contents of the autoclave with agitation attemperatures falling within the range of about 105° C. to about 265° C.at a corresponding pressure of about 20 KPa to about 5000 KPa for aholding time sufficient to cause hydrothermal precipitation, andfiltering the contents of the autoclave to separate the aluminium andfluoride values.